The applicant has developed a hydrometallurgical process for the treatment of copper concentrate so as to produce refined cathode copper by an efficient and environmentally clean process. This process is known as the “CESL™ copper process” and embodiments of the process are described in U.S. Pat. No. 5,645,708 (the '708 patent) the entire contents of which is incorporated herein by reference.
The process as described in the '708 patent in broad terms comprises subjecting the copper concentrate to pressure oxidation in the presence of an acidic chloride solution to produce a solid containing, inter alia, a basic copper salt and a pressure oxidation solution, containing copper in solution depending on the nature of the concentrate and the pH during pressure oxidation. The solid containing the basic copper salt may be subjected to a subsequent acid leaching step, typically at atmospheric pressure, to leach the basic copper salt into solution, thereby obtaining a copper solution which is treated, along with the pressure oxidation solution (should it contain recoverable amounts of copper in solution) to copper solvent extraction and electrowinning to produce cathode copper.
While the CESL copper process as described in the '708 patent, is suitable for the treatment of different copper concentrate compositions, various improvements, modifications or extensions to the process were made to further accommodate different copper concentrate compositions, as well as to provide for the recovery of other metals, such as zinc, cobalt, nickel and precious metals, where these occur in the copper concentrate, or without copper being present in the concentrate. Some of these modifications which pertain particularly to the recovery of precious metals are described in applicant's U.S. Pat. No. 5,902,474 (the '474 patent), the entire contents of which is incorporated herein by reference.
The present invention is concerned with the recovery of precious metals. It is significantly different from the process as described in the '474 patent and has the advantage of being simpler than the earlier process with overall lower costs.
Cyanidation of gold and silver ores is generally carried out in air at ambient pressure. Oxygen is needed in the process, but usually the amount needed is so small that it is adequately supplied by ambient air, which of course contains about 21% oxygen.
Equipment and conditions for cyanidation in commercial plants include:
(a) Agitated (open) tanks, in which the ore is slurried up in a cyanide solution usually with high percentage solids, and leached with free access of air from the surface. The ore must first be crushed and ground to allow it to be suspended by agitation. Several types of agitation have been used including mechanical agitation and air agitation (pachucas).
(b) Retention times are generally 1-3 days. Agitated tanks are used for ores that are relatively high grade in gold or silver, as this method produces better recovery than heap leaching, albeit at higher operating and capital cost. Sparging of air can be done to increase flow of oxygen into the slurry, but that generally increases the loss of expensive hydrogen cyanide into the offgas. Sparging of pure oxygen into the slurry is also known, as is the addition of other reagents such as liquid hydrogen peroxide to accelerate leaching.
(c) Large heaps of crushed ore, which are leached in place by cyanide solution that is sprayed onto or dripped over the surface of the heap. This low cost process is used for low grade ores and takes place over a longer period time of time, generally several months. Air ingress to the heaps by natural processes such as convection is generally sufficient for cyanidation to proceed at an acceptable rate.
With pressure cyanidation of gold and silver ores, oxygen is applied at elevated pressures, i.e. higher than oxygen partial pressure in air at atmospheric pressure, to agitated slurry.
The use of elevated oxygen pressure generally increases the leaching rate of gold and silver. Cyanide leaching is believed to involve oxygen as a reagent, as in the familiar Elsner equation:4Au+8NaCN+O2+2H2O→4NaAu(CN)2+4NaOH
Although the exact stoichiometry of the reaction may be in doubt, (e.g. hydrogen peroxide has been proposed as an intermediate product and reagent), most observers agree that the overall reaction includes oxygen as a reagent.
A similar equation can be written for silver, and with substitution of alternative cyanide reagents, such as KCN instead of NaCN.
Despite several disclosures in the literature of pressure cyanidation processes over many years, commercial application to date has been scarce. Perhaps this is because the obvious advantages of the improved leaching kinetics may be outweighed by the higher capital costs for such a process, with normal gold and silver ores.
However, the inventor is aware of two commercial plants that were built and operated and include pressure cyanidation, namely the Consolidation Murchison plant in South Africa and the Calmet process developed by Calmet of Colorado, USA.
The process employed at Consolidated Murchison was specifically designed to treat some high grade refractory gold concentrates containing arsenic and/or antimony. These rather unusual feed materials required much lower pH conditions than normal, i.e. pH<10, with much higher cyanide concentrations than normal, to obtain satisfactory leaching. Under these low pH conditions cyanide consumption would be very high, using conventional cyanidation (in open tanks) at ambient pressure, because of the excessive volatilization of HCN gas, which increases rapidly with decreasing pH, and also because of the high levels of cyanide needed with this unusual feed material. The process developed at Consolidated Murchison resulted in a commercial plant which satisfactorily resolved these difficulties.
The Consolidation Murchison plant reportedly uses a pipe reactor, operating in batch mode with recirculating slurry, at pH 10 or less, with high cyanide concentrations, and obtains 80-90% gold recovery with acceptable cyanide consumption.
This process is specifically designed for the feed materials on hand at this site, i.e. containing stibnite (Sb sulphide) and arsenopyrite minerals, which requires the low pH, at least low in comparison the usual pH 12 employed elsewhere at this site.
The process of the present invention, as will become apparent below, is capable of dealing with a residue from a copper leach process which has specific components, namely elemental sulphur and cyanide-soluble copper which would otherwise consume large amounts of cyanide with conventional cyanidation processes. In the present invention pressure cyanidation for this feed material is designed to limit thiocyanate production by operating at unusually low retention times. The process of the present invention is not limited to low pH, and the normal pH range is 10-11, but is not limited to this range. The present invention therefore overcomes a different problem than is dealt with by the Consolidated Murchison process.
The Calmet process on the other hand is a combined pressure oxidation and pressure cyanidation process, and was used for a variety of gold containing concentrates including tellurides, sulphides and carbonaceous preg-robbing materials.
Details of the process are scarce, but it seems that pressure oxidation of sulphides in low pH solution was combined with pressure cyanidation of the gold and silver.
It used an agitated autoclave operated in batch mode, in contrast to the pipe reactor at Consolidated Murchison, with a reported capacity of 15-30 tpd concentrate. It was built in Colorado in the 1980's evidently, but apparently closed down a few years later.
The objective of the Calmet process was to treat refractory but high grade gold-silver materials in a one-step pressure oxidation, which simultaneously oxidized sulphides and leached gold and silver. This is different from the present invention, in which pressure oxidation and leaching of base metals, such as copper, if present, is effected first in a separate operation, and the residue from this operation is then subsequently treated for precious metals recovery, as will be described in more detail below.
Gold and silver often occur as trace elements associated with copper in nature, such as in copper sulphide ores. Such ores typically contain 0.3% to 2% Cu, and are usually first subjected to milling and flotation to produce a concentrate of about 30% Cu, which is sufficient to make the subsequent smelting process efficient.
Gold and silver generally follow the copper into the concentrate in fairly high yield, and although they are minor constituents of such copper concentrates, frequently there is enough gold and silver present to be economically significant. Typically, the value of the gold and silver is about 10% of the value of the copper in the concentrate, although this varies widely from one concentrate to another. Rarely is the combined gold and silver content so low as to be negligible.
Occasionally, the combined value of gold and silver in such concentrates is actually higher than the copper, and thus the concentrate is more properly termed a gold or silver concentrate.
When copper sulphide concentrates containing gold and silver are processed by smelting and refining, these metals are generally recovered in high yield from the concentrates (about 90-98%). The extra cost to the copper smelter/refinery of such precious metal recovery is quite low, (incremental to the copper smelter/refining cost itself). The precious metals follow the copper through the various steps of matte smelting and converting to blister copper. The blister copper is then usually refined by electrolysis to remove impurities and during this refining process precious metals report almost quantitatively to the (refinery) anode slime, which has a low mass, (typically only a few kg per tonne of Cu metal). The anode slime therefore has a high concentration of the precious metals relative to the original copper concentrate fed to the smelter, e.g. 1000 to 3000 times more concentrated. Such low mass and high concentrations of gold and silver in the slime lead to low processing costs for final recovery and refining of the precious metals.
The monetary returns from such precious metals in concentrate processed in a smelter are economically significant, and any alternative process for sulphide concentrates (competing with the smelters), must take this into consideration. Smelters will generally pay at least 90% of the value of the gold if there is at least 1 g Au per tonne concentrate. For silver the minimum for payment is about 15-30 g/t. Probably over 85% of copper concentrates traded worldwide have at least this much gold and/or silver, so there is a significant credit for such values when the concentrate payment terms are negotiated between seller and buyer. Typically this credit amounts to about 10% of the value of the concentrate, and generally the gold value is about 80-90% of this, with silver making up the remainder.
Turning to a hydrometallurgical copper recovery process, such as the process described in the '708 patent, if gold and silver are not recovered efficiently along with copper, then the overall economics of the process could be adversely affected in comparison with smelting, even fatally for some concentrates which are particularly rich in gold and silver.
Gold and silver generally do not leach to any significant extent in a hydrometallurgical copper process, and are therefore left almost quantitatively in the residue after recovery of base metals. Therefore, any recovery process for precious metals must be an additional or subsequent step(s) processing such residue, which still has a mass of about 80% of the original concentrate. Concentrations of gold and silver in such residue are only slightly higher therefore in this residue, and still quite low, for example 6 g/t Au and 60 g/t Ag.
In comparison, as mentioned above, anode slimes produced by the smelting and refining process are greatly upgraded from the original concentrate. Thus a typical Cu concentrate with say 5 g/t Au and 50 g/t Ag, might have about 15,000 g/t Au and 150,000 g/t Ag in the slimes that need to be refined.
The challenge to the hydrometallurgical process is the efficiency or economics of the process. Treating such a large mass to recover only small amounts of precious metals at low cost (specifically a cost that is low in comparison to the value of the precious metal content in the residue) is clearly a difficult process to design.
It is possible to leach many gold and silver ores that have very low grades, (even lower than the typical residues of 6 g/t Au and 60 g/t Ag given above), using the well-established cyanide leach process, often with excellent results and low costs. The low concentrations of cyanide necessary to leach gold and silver, and the very low consumption of such cyanide with many such ores, together with the conditions, (ambient temperature and pressure, low corrosion conditions, etc), leads to exceptionally low operating costs for many gold and silver ores, (in terms of $/t ore). This enables ores to be economically treated when the gold content for instance is only 1 g/t Au. Often the cyanide consumptions are less than 0.25 kg NaCN per tonne ore, the cost of which is small compared to the value of gold recovered.
However, the residue generated by a hydrometallurgical process when copper sulphide concentrates are treated has unusual characteristics in regard to cyanidation (compared to naturally occurring ores) which can greatly increase the cost of the process, even to the point of making it uneconomic.
The copper residue contains two components in particular which tend to consume very large amounts of cyanide, i.e. copper and sulphur, when the residue is leached under “standard” cyanide leach conditions, i.e. leaching with dilute(alkaline)sodium cyanide solution at ambient temperature and pressure for 1 to 3 days.
Firstly, the residue still has a significant copper content, despite the fact that it has already been processed specifically for copper extraction. For example, the CESL copper process is about 95-98% efficient for Cu extraction, and thus the residue typically contains 1.2-1.8% Cu. This remaining Cu content is partly (15-25%) soluble in standard cyanide leach conditions, leading to the formation of soluble copper cyanide compounds, such as Na3Cu(CN)4, as well as other cyanide compounds such as sodium cyanate, NaCNO.
Also present in the residue is elemental sulphur which is a by-product of the CESL copper process and typically constitutes 25-35% of the residue. The elemental sulphur also reacts partly with cyanide solutions leading to the formation of thiocyanate compounds, such as NaSCN.
Both of these phenomena lead to very high cyanide consumption with the residue of the CESL copper process, under the conditions of a standard cyanide leach, e.g. 30 kg NaCN consumed per tonne of copper residue, or more than 100 times the consumption typically experienced in leaching gold ores. Such levels of cyanide consumption render the process far too expensive in view of the modest value of gold and silver to be extracted.
At a typical cost of US$ 1.50 per kg NaCN, the cost of cyanide consumed is about $45/tonne of residue. It is worthwhile to consider the effect of such high cyanide consumption on the economics of a typical copper concentrate producing a residue by the CESL copper process.
Assuming a gold and silver content at say 6 g/t Au and 60 g/t Ag, this hypothetical residue has a gross metal value of about $140/tonne of residue, at current prices (time of writing, November 2006).
To the $45/t cyanide cost must be added other reagent costs, (lime and various reagents needed for cyanide destruction), which typically will be at least equal to the cyanide cost alone, leading to a total reagent cost in this case of about $90/tonne residue. Then there are the other necessary operating costs such as labor, energy, maintenance costs plus amortization costs for the capital investment. Typically reagent costs are only a fraction (e.g. 50%) of the total operating cost, so the overall operating cost might be $180/tonne residue, given the original assumption above of 30 kg cyanide/t residue. Thus the total operating costs are of the same order of magnitude as the gross metal value of the gold and silver, with this high cyanide consumption, which clearly renders the process uneconomic.
The total operating cost should not be more than 50% of the value of the metals recovered. Thus to make the process profitable for the example quoted above, (with say 90% recovery of the $140 gross value, i.e. $126), the total operating cost should be no more than $63/tonne residue. Using the assumptions above, the cyanide costs should be no more than 25% of the overall operating cost, or about $16/tonne residue which is equivalent to 10 kg NaCN/tonne residue, in this example.
Thus the typical cyanide consumption for a “standard” cyanide leach on a CESL copper process residue consumes at least 3 times more cyanide than is economically allowable. In addition, the gold and silver themselves cannot easily be extracted from the Cu process residue by cyanide leach solutions.
In summary, with standard cyanide leaching of the copper residue, costs are high, gold and silver recoveries are poor, and the costs of the process tend to outweigh the value of the recovered metals.
It is accordingly the purpose of the present invention to provide a simpler process for gold and silver recovery from sulphide concentrates or other source material which alleviates the above economic challenges.